Method of smelting and refining copper ores and compounds.



R.BAGGALE.Y., METHOD OF SMELTING AND DEFINING COPPER ORES AND COMPOUNDS.

APPLICATION FILED APE. 25, 1907. RENEWED OUT. 22, 1910.

977,996. Patented Dec.6,1$ 10.

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APPLIOATION FILED APR. 25, 1907. RElIEWED OCT. 22, 1910.

977,996, Patented Dec. 6, 1910.

14 SHEETS-SHEET 2.

WITNESSES R. BAGGALBY. METHOD OF SMELTING AND DEFINING COPPER ORES ANDCOMPOUNDS.

APPLICATION FILED APR. 25, 1907. RENEWED 0013.22, 1910.

Patented Dec. 6. 1910.

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R. BAGGALBY. METHOD OF SMELTING AND REFINING COPPER OBIS AND COMPOUNDS.

APPLIOATION FILED APR. 25, 1907. RENEWED OUT. 22, 1910.

977,996. Patented Dec. 6, 1910.

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R.BAGGALBY.

METHOD OF SMELTING AND REFINING COPPER ORFS AND COMPOUNDS.

APPLIOATION PILED APR. 25, 1907. mmwsn OUT. 22, 1910.

977,996., Patented Dec.6,1910.

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R. BAGGALEY.. A METHOD OF SMELTING AND REFINING COPPER OBI-IS ANDCOMPOUNDS.

APPLICATION FILED An. 25, 1907. RENEWED 001.22, 1910.

Patented Dec. 6,1910

14 SHEETS-SHEET a.

INVENT R' 41444 49 "m: NORRIS PETERS ca., WASHINGTON, n4 cv R. BAGGALBY.

METHOD OF SMELTING AND REFINING COPPER ORES AND COMPOUNDS.

Patented Dec. 6, 1910.

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R. BAGGALEY. METHOD OF SMELTING AND REPINING COPPER ORES ANDGOMPOUlfIDS.

APPLIUATION FILED APR. 25, 1907. RENEWED 001'. 22, 1910.

977,996. Patented Dec. 6, 1910.

14 SHEETS-SHEET 8.

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METHOD OF SMELTING AND REFINING COPPER ORES AND COMPOUNDS.

APPLIOATION FILED APR. 25, 1907. RENEWED 001. 22, 1910.

977,996. Patented Dec. 6, 1910.

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I R. BAGGALBY. METHOD OF SMELTING AND REFINING GOPIER ORES ANDCOMPOUNDS.

APPLICATION FILED APR. 26, 1907 RENEWED OUT. 22, 1910.

977,996. Patented Dec.6,1910.

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R. BAGGALEYJ METHOD OF SMBLTING AND REFINING COPPER ORES AND COMPOUNDS.

APPLICATION FILED APR. 25, 1907. RENEWED 001222, 1910.

977,996. Patented Dec. 6, 1910.

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METHOD OF SMELTING AND REFINING OOPPLR OBES, AND GOMPOUNDS. APPLICATIONFILED APR. 25, 1907. RENEWED 001222, 1910. 977,996.

INVENTOR nu: NORRIS PETERS co., WASHINGTON, n. c.

WITNESSES R. BAGGALEY. METHOD OF SMBLTING AND RBFINING COPPER ORES ANDCOMPOUNDS.

APPLICATION FILED APR. 26, 1907. RENEWED OUT. 22, 1910.

977,996. Patented Dec. 6, 1910.

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R. BAGGALEY. METHOD OF SMELTING AND REFINING COPPER 0112s AND COMPOUNDS.APi'LIOATION FILED APR. 25, 1907. RENEWED OUT. 22, 1910.

977,996. Patented Dec. 6, 1910.

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: 1 Q i g i 5 i 3 I 59 1 WITNESSES INVENTOR NORRIS PETERS co,WAsnIncrmJ. 0 c4 RALPH BAGG-ALEY, OF PITTSBURG, PENNSYLVANIA.

METHOD OF SMELTING AND REFINING COPPER ORES AND COMPOUNDS.

Specification of Letters Patent.

Patented Dec. 6, 1910.

Application filed April 25, 1907, Serial No. 370,238. Renewed October22, 1910. Serial No. 588,535.

To all whom it may concern:

Be it known that I, RALPH BAGGALEY, of Pittsburg, Allegheny county,Pennsylvania, have invented a new and useful Method of Smelting andRefining Copper Ores and Compounds, of which the following is a full,clear, and exact description, reference being had to the accompanyingdrawii'igs, in which- Figure 1 represents a longitudinal crosssection ofmy primary converter for eliminating metalloids, performing in thisrespect the function of a blast furnace for the treatment, withoutcarbonaceous fuel, of such ores as contain within themselves sufficientsulfur or other oxidizable elements and compounds, such as arsenids,antimonids, sulfids, &c., to produce the heat with which to conduct acontinuous smelting operation. Fig. 1 also illustrates in longitudinalcross-section, a forehearth and inclosed drop tube, which exclude theoutside air and thus prevent chilling the molten slag and matte; Fig. 2is a cross-section on the line IIII of F ig, 1; Fig. 3 illustrates acrosssection on the line III-HI of Fig. 1; Fig. t is an alternate formof my primary smelting converter, together with its forehearth. Thisalternate form may be used when it is preferred that such primaryconverter shall remain in a fixed position, or in other words, when itis not desired to rotate the vessel for the purpose of emptying thecontents, should a stoppage from any cause become necessary. It is alsouseful in treating very refractory ores, which may demand a very longtravel, while submerged in the molten bath of matte to dissolve them;Fig. 5 represents a cross-section on the line 55 of Fig. 4:; Fig. 6represents a side elevation of my rotatable primary smelting converterwith its inclosed drop tube, for-hearth and waste slag-car, incombination with one form of a preheater. The latter is intended toexpel moisture from the ore-feed and to preheat the ores to a degreeless than the fusion point of the contained sulfid, arsenids,antimonids, &c., by means of the waste hot gases that constantly escapefrom the nose of the vessel while ores are being smelted in it; Fig. 7represents a cross-section of Fig. 6. It illustrates the method ofdrying and preheating the ore, or if preferred, of discharging the hotgases emanating from the process, directly into the dust chambers, flueand stack, without first passing the same through the preheater; Fig. 8illustrates a longitudinal cross-section of my secondary or ironeliminating converter. The vessel is shown with a water-cooled, solidmetal top or throat, in order to illustrate various forms ofconstruction; Fig. 9 represents a cross-section of Fig. 8; Fig. 10illustrates a longitudinal cross-section of my preferred form of copperrefining furnace; Fig. 11 illustrates a cross-section on the line X1, XIof Fig. 10; Fig. 12 illustrates a front elevation of Fig. 10, showingmeans for injecting saw-dust or other gas producing media; Fig. 13illustrates an end view of Fig. 12; Fig. 1d illustrates my preferredmethod of recovering mineral values from fines and fluedust, by minglingthese with low grade molten matte before the same is placed in thesecondary converter for bessemerizing.

I have considered it unnecessary to include in these illustrations acopper blast furnace or cupola for the melting of ores containing lessthan 56 per cent. of oxidizable elements and compounds, or in otherwords, for the first fusion of such ores as demand that certainpercentages of coke be added to supply the deficiency of natural fuelscontained in the ores themselves, in order to successfully subject themto a simple melting for the elimination of the silica, alumina, lime orother metalloids. Any of the ordinary forms of copper blast furnace orcupola in general use may be utilized for this purpose.

Ores as usually found in nature contain an enormous proportion of wasteand a comparatively minute percentage of mineral. A 2% copper ore, forinstance, which is about the present limit of possible profitableextraction, when the values are not in the form of virgin copper,contains 40 pounds of valuable mineral and 1960 pounds of waste per ton.The problem is to separate the one from the other quickly, cheaply andwith minimum losses. All processes of separation, whether by water, oil,air or smelting, are attempts to do this.

When highly silicious Butte ores are smelted they demand the addition oflarge percentages of non-mineral bearing fluxes principally in the formof lime and iron, to make this separation possible. Such additions havethe effect of greatly increasing the consumption of coke and they doublethe volume-hence the mineral losses in the slag. hen such ores arecrushed and concentrated by water, oil or air, the mineral losses areexcessive. one-third of the total mineral contents.

All present methods of extracting the mineral values from copper, goldand silver ores include some one of these various forms ofconcentration; such, for instance, as water concentration, or where theores cannot be crushed and concentrated, owing to their physicalcondition, they are usually roasted in open roast heaps or in stalls,and they are thereafter subjected to smelting in blast furnaces withheavy percentages of lime and colre in order to produce what is calledin the trade it converter matte; or in other words, a matte containingabout 50 per cent. of copper. Such matte is thereafter bessemerized invessels containing heavy bodies of crushed silica, in the form of alining, which, through its corrosion and destruction, is depended uponto unite with the iron, after oxidation, in the matte and thus pass offin the form of silicate of iron slags.

All present processes of recovery are expensive and are extremelywasteful of the minerals contained in the ore. The records for manyyears past in the Lake Superior district show that in the waterconcentration process alone, ores containing primarily 3 per cent. ofcopper, only yield, after water concentration, 2 per cent. In otherwords, one-third of the entire mineral contents is lost in this singleprocess. In the Butte district of Montana, the loss of mineral hasaveraged more than one-third.

The object of my present invention is to provide an entirely new methodof extracting the mineral values of copper, gold and silver ores a.lVithout the cost of a watersupply and of a water concentration plant.Z). W'ithout the enormous percentages of mineral loss that now prevailin all water concentration and other processes. 0. ithout the mineralloss that experts know re sults from all forms of heap or stall or otherroasting. (Z. Without the cost of coke or other carbonaceous fuel insome cases, and with less than one-half present costs, in others. 6.With less than one-fifth the slag loss that prevails in any present formof smelter practice. f. lVithout the present excessive mineral loss thatresults from saturation in furnace bottoms and side-walls. 9. Withoutthe fuel and labor costs of remelting an excessive proportion ofcongealed converter slag and smelter scrap. h. Without creating thepresent poisonous tailings nuisance, that has been found to be soinjurious to land and to water in the vicinity of ore treatment plants.9'. lVithout creat ing more than a small proportion of the presentsulfur and arsenic fumes nuisance, that is so objectionable toagricultural interests for many miles around all copper smelting plants.lVithout requiring more than live hours in which to produce good Thesegenerally amount to blister copper out of ores, where hours is now theaverage in present practice. h. Without requiring an investment of morethan one-fifth that demanded by all present practice in treatii'ig thesame tonnage. Z. 'lhat renders available and commercially valuable largebodies of low grade ores that are relatively low in copper or inprecious metals, yet which contain ample percentages of natural fuel,principally in the form of sulfid, for their own refining. Enormoustonnages of such ores exist on the island of lk'ewfoundlaml andelsewhere, yet these great deposits are not and cannot be cominerciallyworked by any present process, ex" cepting that described in my presentinvention.

The enormous advantage of this process, when viewed as a whole, over anyexisting process, to accomplish the same results of extracting copper,gold and silver from their ores may be best shown by the approximatecost of each.

The cost of producing copper by all pres ent processes may safely besaid to average more than 9 cents per pound. YVith my present process itis safe to say that in treating ores that contain the requisitepercentages of natural fuels, that is to say, oxidizabl-e elements andcompounds in excess of 56 per cent, that copper can be thus produced atone-half the cost of present practice at all the present plants,whatever the same may on test prove to besat each. "With ores that aredeficient in natural fuels and which contain less than 56 per cent. oftotal oxidizable elements and compounds, so that coke or othercarbonaceous fuel must be added to a greater or less degree, governed bythe amount of natural fuel in the orcs themselves, and which additionsmust be made of carbonaceous fuel in order to main tain a successfulsmelting process, then the difference in cost is not and cannot be sogreat, because in this first fusion a blast furnace must be used, withat least some of its wastefulness and its objectionable features. Inthis event, however, it is safe to state that copper by my process canthen be produced at two-thirds the cost of present practice, or forless, according to the percentages of coke that must be added to theores, in order to supply the deficiency in their natural fuels.

In treating ores by this process they must of necessity be divided intotwo great and radically different classes, to wit:

First. Those that contain sutlicient natural fuel within themselves tomake their treatment by smelting with it alone possible. In this classare included all those that contain in excess of say 56 per cent. oftotal oxidizable elements and compounds such as sulfur, iron in thesulfur form, arsenids, antimonids, etc. hen a blast of air is driveninto any or all of these while in the molten form, heat results fromtheir transition from the primary into the ulti mate or oxid form.Everything on the crust of the earth that is not already an oxid isundergoing continuous oxidation through the processes of nature andthrough the action of the oxygen contained in the air and in themeteoric or surface waters. If a piece of steel or iron, although it isalready an oxid in one of its forms, be exposed to the weather, it maydemand ten years to be transformed into rust, or in other words into itsfinal oxid form. This transition produces heat to the same degree thatit is produced if the same piece of iron or steel were oxidized withgreat rapidity. In the one case the oxidizing process is so slow thatthe heat generated is imperceptible. In the other it may be so rapid asto produce silver white incandescence. The ores referred to atNewfoundland; those found at Rio Tinto, in Spain; those at Tharsis, inPortugal; those at Mount Lyell, Tasmania, as well as many otherimportant deposits in the world are all included in this class. Some ofthem contain as high as per cent. or even more of total oxidizableelements and compounds. All such are capable of successful treatment bysmelting, through the heats of oxidation alone, in my primary smeltingor metalloid eliminating converter, and entirely without the addition ofcarbonaceous fuel of any lrind. he efficiency of this primary converterand its smelting capacity will be materially increased by utilizing thehot gases emanating from the process, in drying and in preheating theores before they are fed into the vessel. This preheating is notessential to success, from a purely metallurgical standpoint, as rawores direct from the mines may be successfully smelted thus, withoutpreheating. The efiect of preheating is simply to increase theefficiency of the apparatus and to materially augment the tonnage of orethan can be dissolved or melted in it in a given time. For ores of thisclass, my preferred method of treating them in my primary smelting ormetalloid eliminating converter, with its forehearth, may be describedas follows: A bath of clean matte is used with which to start theprocess. It may be derived from melting ores in a cheap cupola andseparator or from any other source. It is made preferably from a simplesmelting (without any attempt at concentration) of such ores as are richin sulfur, iron, &c. hat are called Butte smelting ores, which contain 6per cent. or more of copper, may be used because these are comparativelylow in silica and high in various oxidizable elements and compounds.Concentrates, pyrrhotites and many sulfide, arsenids, and antimonids aresuitable for the purpose. After the process has been started and theoxidizing blast has been turned on, the iron supply of the bath isthereafter continued indefinitely by feeding continuous charges of rawores of this description. The character of the bath may be changed orregulated at will, by the kind of ores fed. Fuel ores will increase theheats and the iron contents. Silicious ores will flux off and reduce thepercentage of iron and they will usually increase the valuable mineralcontents. The latter will be governed by the contained values.

The primary converter is the equivalent of and is designed to supersede,while treating suitable ores, the ordinary copper blast furnace, and itsore charges are intended to correspond, the difference being that itdoes not require coke and non-mineral bearing fluxes and it does notcreate a reducing action during fusion. At the same time it will berecognized that its limitations must always be to such ores as containwithin themselves su'liicient natural fuel to make a continuous smeltingprocess possible by the heats of oxidation alone.

The more basic the bath, the larger the proportion of highly siliciousores will it be possible to dissolve in it. After the primary converterhas been thus charged the converting blast is turned on in great volumeand the ore-feed is commenced. The latter may be continuous orintermittent, as preferred. The vessel is preferably 20 feet in lengthand 8 feet in diameter, and it is provided with say twenty-eight 1?,inch twyers, if of the rotatable type, as illustrated in Figs. 1, 2 and3. If it be of the fixed or alternate type, as illustrated in thedrawings, it may be of greater length and the number of twyers should beincreased according to its length.

Before entering the vessel, the ores are preferably passed through apreheater of some form; such, for instance, as that shown in thedrawings, and the hot gases which continuously escape in enormous volumeare utilized to preheat them, without cost for fuel, to a degree lessthan the fusion point of sulfids, arsenids or antimonids, or say, forinstance, up to 700 degrees F. The heat may be utilized to the fullestextent without any cooling influence from an adn'iixture of outside air,because this vessel remains continuously in its upright position so longas work in it progresses. This pre-heating, as before stated, has theeffect of expelling all moisture and it heats the ores to an importantextent. Each degree of heat thus added, without cost, materiallyaugments the smelting capacity of the vessel in a given time. The orecharges enter the vessel at one end and they float slowly toward theopposite end while sub ected to the ntensely corrosive action of a lowgrade, basic matte and to the intense heats produced by an enormous airblast acting upon a low grade bath. The ores are thus quickly dissolvedor melted, without carbonaceous fuel. Inasmuch as they are thus droppedinstantly into the smelting temperature which always exists in the bathand above it in the vessel and which must always exist so long as thebath remains molten, the sulfids, arsenids, antimonids and other fuelscontained in the ores are liberated and are forced to oin the matte,without volatilization; hence they become available as heat producers incon tiuuing the smelting process. Careful tests have proved that acontinuous temperaturein excess of 2000 degrees F. is demanded in orderto maintain a bath of 10 per cent. copper matte in a molten, or in otherwords, in a liquid condition.

A great excess of silicious ore may be fed into this primary smeltingconverter, if desired, beyond the proportion that can be fused and thatwill form selective slags. The mineral values contained in these areextracted by liquation and they join the molten bath of matte. The cleansilica shells float away and pass off with the worthless slag. Suchseparation by liquation is found to be exceptionally clean. Any occludedparticles of mineral will be fused and join the molten bath at or about700 degrees 0. During their sojourn in the primary smelting converter,ores are continually subjected to ten'iperatures in excess of 1300degrees C. The silica will fuse and form selective silicate of ironslags in suitable proportion to the iron that exists in the bath. Anyexcess will remain unfused and will yield up its values throughliquation alone, as stated.

Should it be found in practice that slags become sticky or pasty, eitherfrom a temporary excess of heat sufficient to cause a ten'iporary excessof molten silica in the bath, from an excess of Zinc, or from any othercause, limestone may temporarily or continuously be fed either into theprimary or into the secondary converter. I much prefer, for thispurpose, the temporary feeding of fluorspar; because T have found inpractice that a relatively small proportion of it will at once liquefythe slag and remove any tendency to pastiness to a most astonishingdegree. I would also call attention to the fact that in the primarysmelting converter, l have found that Zinc does not create thick, pastyslags, as is so common in blast furnace work. In the latter it has beenfound that 5 or 7 per cent. of zinc represents about the maximum limitthat can be successfully worked, whereas in the smelting converteii',ores containing 20 per cent. can be easily and successfully treated. Thereason for this I have found to be that in the smelting converter muchthe greater proportion of the zinc is oxidized or volatilized and passesoff in the form of hot gases, so that only a small proportion of itremains in the slag. This explanation will account for the greatdifference in percentages that may be successfully worked in the twoprocesses of fusion.

The primary converter, at the end opposite to the orefeed, provided witha water-jacketed, untrapped overflow-spent in its axis. its the smeltingprogresses, matte, slag and nnfused silica ore shells constantlyoverflow, the same as from a copper blast furnace. These together dropinto a forehearth, or v cparuting vessel that is prelimably 20 fee inlength and say 10 feet in width, where an exceptionally clean separationoccurs. The matte constantly accumulates at. the bottom of theforehearth. It is not desired that this shall exceed a grade of about 15per cent, in ordinary practice. It may be raised or lowered to harmonizewith special conditions.

The worthless slag overflows quickly and continuously from the converterinto the forehearth and from the forchearth it over flows contimiouslyand passes direct to the dump. l t essential to success in this form ofsmelting ores by dissolving them in matte that the surface of the mattebody he kept at all times clean and free from floating slag, so that theentering ore feed can drop into and instai'itly become sul'in'ierged inclean matte. In this apparatus the over flow-spout is always open, freeand untraoped. No floatin bod of slap can. i :1 Q ther fore, accumulateas it does in a blast furnace with a trapped spout. Un the contrary, thefloating slag passes ojll' quickly and the matte lath remains clean. inthe forehearth the slag is always discharged from a body of very lowgrade matte.

will be noted that in this primary or metalloid eliminating converterpractically all of the silica, alumina, lime and other metalloids are atonce separated andv passed to the slag dump, with absolutely minimummineral losses. The product consists of practically pure iron, copper,gold, silver and sulfur. in other words, the metals that it desired torecover remain associated with a certain percentage of iron and sulfur,which together usually approximate 15 per cent. of the total tonnage ofore, while at this one point 85 per cent. of the total ore tonnage isusually at once discharged to the dump, at a minimum cost and withminimum mineral losses.

Particular attention is called to the fact that this process eliminatesthe poisonous tailings nuisance, that has been the cause of dan'iage inagricultural districts. The water concentration process not used and theslugs produced are not objectionable. A more important item, however, isthe mitigation of the nuisance of sulfur and arsenic fumes by thisprocess. When ores containing sulfur and arsenic are fed into a blastfurnace and travel slowly downward from the level of the charging-floorto the twyer zone, or in other words, to the only point in the blastfurnace where it is possible to melt them, they become gradually heated.The distance varies in different forms of furnace, say from 9 to 16feet, and during all of this time the sulfur and arsenic are beingexpelled, at least one-half of which is in the free state, so that theypass out of the stack in the form of gases 80,, SO, and are thenprecipitated upon agricultural districts for twenty miles around thesmelter plant. Inasmuch as the ordinary copper smelter plant thus ejectsfor subsequent deposition on the farm lands of the district from 200 to600 tons of sulfur every twenty-four hours, according to the smeltingcapacity, the magnitude of this offense will be understood andreasonable people cannot find much fault with ranchers and farmers forpursuing litigation and for obtaining injunctions, as has been the casewithin the past two or three years. The wonder is, to disinterestedpeople, that more injunctions have not already been obtained. The sulfuris volat-ilized and passes out of the stack in the form of gases, assoon as the ores reach a temperature in the blast furnace of about 700degrees C. The ores cannot be melted, so that the matte can absorb thesulfur until a temperature of more than 1190 degrees C. has beenreached. This temperature cannot be produced in a blast furnaceexcepting only in the twyer zone. Long before the ores reach this pointin the furnace most of the sulfur has passed out of the stack in theform of In my method of smelting ores containing very high percentagesof sulfur, in the pri mary converter, the fact is emphasized that allthe heats utilized in the smelting process are obtained solely from theabsolute combustion of the sulfur and of all other oxidizable elementsand compounds contained in the ores or in the matte. This combustionoccurs within the vessel itself, otherwise the heats would be uselessfor smelting purposes. The heats that are produced and that arecontinuously maintained in this process of smelting are at a silverwhite incandescence and represent about 2300 degrees F. Owing to thesehigh heats, the combustion within the vessel at all times very thorough.

Second. Those ores that contain less than 56 per cent of snlfids,arsenids, and antimonids or oxidizable elements and compounds of anykind; hence these that demand an addition of carbonaceous fuel, as ameans of providing sufficient heat with which to smelt them.

It is not pretended by me that I am able to smelt ores without fuel ofsome kind; but where the ores within themselves contain suflicientnatural fuels, I utilize these in. this process and I have found thatunder these conditions it is unnecessary to add carbonaceous fuel of anykind in order to maintain indefinitely a successful smelting process.lVhere ores do not contain suflicient natural fuel within themselveswith which to smelt them, I then supply only enough carbonaceous fuel tomake up such deficiency. This added fuel cannot be handled to advantagein my primary or smelting converter, and as aconsequence, in the firstfusion with all ores that contain less than 56 percent. of totaloxidizable elements and compounds, I am compelled to utilize a blastfurnace, excepting with say 10 or 15 per cent. of such ores, as will behereinafter described and which are dissolved in the secondaryconverter. My method in this process of treating ores in the blastfurnace is radically different from that which has prevailed in all pastpractice. I add only enough coke to make up the deficiency of naturalfuels, as stated, and the sole object of this first fusion is for thepurpose of eliminating the silica, lime, alumina or other metalloids. Noattempt whatever is made at concentration, beyond what naturally resultsfrom a simple melting of the ores, to accomplish the elimination of themetalloids alone, as stated.

Where, in present and past blast furnace practice, 18 or 20 per cent. ofcoke has been found necessary in treating Butte and many other ores,supplemented by water concentration and roasting in heaps, in stalls orroasters, in order to produce a 50 per cent. converter matte, I avoidthe expenses and the mineral losses of all crushing, all forms ofconcentration and roasting, as well as the use of about one-half of thecoke in the blast furnace. In lieu of all this, I subject about 90 percent. of the entire ore tonnage produced from the mines, of whatevergrade it may prove to be, and usually not exceeding per cent. of copper,to a simple melting in the blast furnace. I use for this purpose only 7or 8 per cent. of coke and never more than 10 per cent; or in otherwords, less than one-half the amount that is now used in generalpractice. The percentage of coke thus used in my process will begoverned by the percentages of natural fuels contained in the ores undertreatment. This first fusion or simple meltwithout any attempt atconcentration) ing results in the accun'iulation in the fore hearth of abody of matte that usually varies from 15 to 25 per cent. in copper.This percentage, of course, will be governed absolutely by the mineralcontents of the ores. I pay no regard to the mineral contents of From apurely metallurgical the matte.

in this process? standpoint it may be anything from 3 per cent. to 30per cent. The richer the ores under treatment and the higher as aconsequence the mineral contents of the resultant matte, the better willbe the commercial results. The point that I wish to particularlyemphasize, as a means of showing that this process of smelting in ablast furnace is radically and fundan'ientally different from all pastpractice, and that it is in effect an absolute reversal of all pastpractice, is that in this first fusion 1 do not attempt or desire anyconcentration beyond what results from a simple melting of the ores,solely as a means of eliminating the worth less metalloids. In all othersmelting prac tice, concentration is desired and is attempted to theutmost degree, at each and every step. The question naturally arises:hen others desire such concentration in this first fusion, why do I seekto avoid it The answer is simple: Concentration means an increased lossof iron in the matte. I wish to retain in my matte every unit of ironthat it is possible for me to retain, because each unit of iron thussaved means an increased amount of mineral bearing, arofitable ore thatcan be melted, without cost, in the secondary stage of this process.This, then, is one object in producing a low grade matte, or in otherwords, a matte that contains heavy percentages of iron, which I havefound in practice usually is about 67 per cent. As before stated, thesole object of this first fusion, whether the same be accomplished inthe primary or smelting converter, while treating ores that containample percentages of natural fuel for their own reduction, or whethertreating in a blast furnace ores that are so deficient in natural fuelsthat this deficiency must be made up in the form of coke, is toeliminate the silica, alumina, lime and other worthless metalloids. Ihave not considered it necessary to include in the drawings apparatusused in practicing this portion of the process, because any of the usualforms of blast furnace or cupola are suitable.

Either one or the other of the foregoing simple fusions, while treatingores of widely different character in radically different apparatus andat widely different costs, will result in approximately the sameproduct. This is low grade matte, amounting generally to about 15 percent. of the total ore tonnage, and it consists of the copper, gold andsilver contained in the ores, together with a very large proportion ofthe total iron contents and with a large proportion of the sulfur.

It will be noted that approximately 85 per cent. of the total oretonnage, or in other words, the metalloid contents of the ores undertreatment, has been discharged to the dump, from the respectiveforehearths and from a body of very low. grade matte, that rarelyexceeds 20 per cent. of copper and as a rule is much less. in acontinuous eight months run of the Pittsmont smelter at Butte, I foundthat the mineral contents of this slag was usually the one-twentieth ofone per cent. in copper. Sometimes it ran so low as to make itimpossible to analyze by any of the ordinary methods. it never exceededthe one-tenth of one per cent. throughout our entire run. The average ofall present practice may safely be stated to show a mineral loss in theslag of three-fourths of one per cent. .lt never averages less thanfortyfive lnnidredths of one per cent, and a large proportion of it isnow running from 1 to 1-3 per cent. Some of the old slag dumps at Butte,in Montana, are now being blasted out and mined for shipment to theWashoe smelter, for use as an iron flux, because oxidized iron veinoutcrops are extremely scarce in that section. These slags contain froml-lto 3-1, per cent. of copper and they illustrate past treatmentlosses, which show nearly a much copper in the slags as in the ores.

When it is remembered that in Butte practice approximately 50 per cent.of flux is used for the privilege of smelting 50 per cent. of mineralbearing ores, and that this flux makes slag in volume equal to that madeby the ore, and that all such slag carries away to the waste dump anequal percentage of copper, it will be realized that present practice atButte means a slag loss of nine-tenths of one per cent. of copper, ornearly one-third of the total copper contents of the ore, because thetotal ore product at Butte has averaged for many years past 3} per cent.in copper, and it is to-day said to average 3 per cent. in copper, forthe reason that the mines are at present working on much lower andricher levels than has ever been the case in the past. hile thisworthless slag overflows from the forehearth continuously to the wastedump, the low grade matte product accumulates in the forehearth ofeither form of apparatus, and it is from time to time tapped out, forfurther treatment in the secondary stage of my complete process, or inother words, in my secondary converter, for the elin'iination of itscontained iron, on lines that result in profits, where all past andpresent practice outside of this method of treating ores, constituteswhat is well known in the trade to be about the heaviest single item ofexpense in the cost of producing copper.

Recooery of mineral values from flucal'usz' muZ flacs.Onc of the manycomplex problems in present treatment processes is the recovery ofmineral values from flue-dust and fines. Briqueting is usually resortedto, in an attempt to recover these values, but

